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CN103290224B - Recovery process for valuable metals in tungsten residues - Google Patents

Recovery process for valuable metals in tungsten residues Download PDF

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CN103290224B
CN103290224B CN201310210077.7A CN201310210077A CN103290224B CN 103290224 B CN103290224 B CN 103290224B CN 201310210077 A CN201310210077 A CN 201310210077A CN 103290224 B CN103290224 B CN 103290224B
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CN103290224A (en
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刘甲祥
徐雪明
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Beijing Hengtai Yanlei Technology Co ltd
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Abstract

The invention provides a recovery process for valuable metals in tungsten residues. The tungsten residue treatment method comprises the following steps of: boiling tungsten residues and 15%-25% of hydrochloric acid for acid dissolution, further adding NaF and sodium nitrate, and filtering to obtain filtrate after the end of the reaction; then extracting tungsten from the filtrate by adopting N235, and evaporating, crystallizing and drying a stripping solution to obtain APT (ammonium paratungstate); and performing alkali dissolution on filtration residues, and performing pressure filtration to directly recover Ta, Nb and Sn valuable metals. Tungsten, scandium, iron, manganese, tin, copper, tantalum and niobium valuable elements and the like in the tungsten residues can be comprehensively recovered, thus the recovery cost is reduced. The recovery process disclosed by the invention further has the advantages of good environmental benefits and low cost. According to the recovery process disclosed by the invention, the acid is adopted for dissolving the tungsten residues, the tungsten and the scandium are recovered under acid conditions, the other elements are simultaneously recovered, and the process is simple.

Description

The recovery process of valuable metal in tungsten slag
Technical field
The present invention relates to the recovery process of valuable metal in a kind of tungsten slag.
Background technology
Wolframite gets the residue of gained after sodium wolframate through alkali lixiviate, is usually referred to as tungsten slag.Produce domestic accumulation decades to have reached more than 1,000,000 tons and do not deal with, and also have new tungsten slag to produce every year.These slags not only cause very large burden to environment, and waste the nonferrous metal resource of country to a great extent.Because in tungsten slag except tungstenic, also containing the non-ferrous metal (each constituent content of tungsten slag is in table one) such as iron, manganese, scandium, molybdenum, tantalum, niobium, it is a very abundant nonferrous metal resource.From tungsten slag, reclaim non-ferrous metal, domestic and international existing operational path is mainly summarised as wet processing and pyrogenic process-wet method and combines process.
Wet processing is the method that current most countries adopts, because tungsten slag is the important source material carrying scandium, so it is many to propose the research of scandium to tungsten slag.Hunan Research Institute of Rare Earth Metal Materials just adopts sulfuric acid to leach from the eighties in last century, P204+TBP+ kerosene extraction, NaOH back extraction, dissolving with hydrochloric acid, ammonia precipitation process, and the processes such as oxalic acid precipitation, successfully prepare High-purity Sc Oxide.What Zhuzhou Hard Alloy Plant adopted carries scandium from tungsten slag, is also to adopt sulfuric acid leaching, with P204+TBP+ kerosene extraction scandium, and NaOH back extraction, dissolving with hydrochloric acid, Na 2hPO 4except Ca, mandelic acid, except Zr, PMBP+ benzene is except rare earth, so repeatedly several times, obtains the Scium trioxide of 99.999%.What Xu Tinghua etc. proposed carries scandium from tungsten slag is also adopt sulfuric acid leaching, and P507 extracts, H 2sO 4+ H 2o 2+ HF back extraction.The molten material of Liu Huizhong proposition sulfuric acid, N1923 extracts, hydrochloric acid back extraction, oxalic acid precipitation, obtains thick Scium trioxide after calcination.Zhong Xueming also proposes with sulfuric acid leaching, and the N1923 removal of impurities of different concns, scandium-enriched, ammonia precipitation process, dissolving with hydrochloric acid, obtains Scium trioxide after oxalic acid precipitation.Obviously, wet processing can not do synthetical recovery to the valuable element in tungsten slag, cause the wasting of resources, and tooling cost is higher.
The facture application that combines of pyrogenic process-wet method is less, and main flow is that fire reduction smelts ferrimanganic tungsten, uses concentrated sulfuric acid dissolution smelting slag, passes through P 204, TBP, sec-octyl alcohol and kerosene composition organic extractant phase, drench chromatograph and naphthenic acid by TBP extraction again after back extraction and purify, after oxalic acid precipitation, calcination becomes 99.999% Scium trioxide.This method is than more comprehensive synthetical recovery, but pyrometallurgical smelting energy consumption is high, and environmental pollution is large, and with the energy-saving and emission-reduction advocated now, the developing direction that cleaning ambient type is produced not too conforms to.
Summary of the invention
In order to solve the problems of the technologies described above, the invention provides one can carry out synthetical recovery, low cost to tungsten, scandium, iron, manganese, tin, copper, tantalum niobium valuable element etc. in tungsten slag, and has the recovery process of valuable metal in the tungsten slag of good environmental benefit.
The present invention includes following steps:
1) concentration first adding calculated amount in cell body is 15 ~ 25% hydrochloric acid, be heated to 60 DEG C, black and white mixing tungsten slag is added at leisure under whipped state, after adding black and white mixing tungsten slag, add 15 ~ 25% hydrochloric acid of 0.2 times of first time hydrochloric acid content again, after heated and boiled, add the NaF of 20kg/t, the NaNO of 10kg/t 3; When pH value to 1.0 time, then add 15 ~ 25% hydrochloric acid of first time hydrochloric acid content 0.4 ~ 0.5 times amount, make S:L=1:4.5, boil when becoming reddish-brown to mud and stop heating, continue to stir, add water 2t/t, stirs 0.5h;
2) when step 1) below the feed temperature to 80 prepared DEG C time, carries out filter press, after press filtration, water washing is added to filter residue, and air blast dries up; Filtrate is put into storage tank and is left standstill at least 12h, then carries out essence filtration;
3) first toward groove in, add water, then by the 2nd) step press filtration wash after filter residue add, add 50 ~ 80kg/t NaOH simultaneously, add the NaF of 20mg/t, the Na of 20mg/t 3pO 4, L/S=1.5, heated and boiled 30 minutes, be incubated 1 hour, control endpoint pH is 10-11;
4) when alkaline solution temperature is down to below 80 DEG C, filter press is carried out, and with the washing filter residue 1 ~ 2 time of 30-40 DEG C; Slag rate 5-10%, Ta, Nb, Sn enrichment 8-16 doubly, can directly reclaim Ta, Nb, Sn valuable metal;
5) stirring and be 10% hydrochloric acid neutralization procedure 4 by concentration under boiling conditions) filtrate that produces to NaOH4-5g/L, add NH after boiling 30 minutes 4cl solution to pH value is 8 ~ 9, and sedimentation and filtration is except white residue;
6) adopt strong basicity gel-type anion resin by the filtrate ion-exchange absorption tungsten after molten for alkali silica removal; Wire velocity control 6-10cm/min during exchange, WO in liquid after inspection exit ion-exchange 3content; Leacheate: NaCl 4-8g/L or NaOH10g/L, drip washing linear velocity 2-4cm/min, stripping liquid is: 3.5mol/LNH 4cl or 2mol/L NH 3h 2o, desorb linear velocity: 24cm/min, stripping liquid consumption is 50% of resin volume;
7) by step 2) after the filtrate reduction of filtering of the essence that produces, the organic phase of 10 ~ 30%N235,10 ~ 20% sec-octyl alcohol and kerosene composition, squeeze in extraction tank by comparing O/A=1/20-30, stir 2 minutes, 4 stage countercurrent extractions, then add water by O/A=1/1, stir 3 minutes, 6 grades of washings; Then add the NaOH solution of the 2mol/L of heat by O/A=2/1, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then add water backwash by O/A=1/3 ~ 5, stir 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then add the hcl acidifying of 6mol/L by O/A=3 ~ 5/1, stir 3 ~ 5 minutes;
8) squeeze in enamel reaction still by stripping liquid, calorify 80 ~ 85 DEG C of evaporative crystallizations, endpoint pH is 7 ~ 8, crystallization control rate 90% ~ 95%, crystal vacuum filtration, and with 30 ~ 40 DEG C of water washings, washing lotion and mother liquor return absorption;
9) by the crystal that washed 90 ~ 120 DEG C of dryings, obtain ammonium paratungstate.
In above-mentioned tungsten slag valuable metal recovery process in, also comprise Sc 2o 3preparation process:
1)) by extraction W after more than liquid squeeze in extraction tank, add the organic phase of 12%P204,4%TBP and 84% kerosene composition, compare O/A=1/20, stir 3 minutes, 4 stage countercurrent extractions, control Scium trioxide content in raffinate and be less than 3mg/L, add that concentration is 15%HCl, concentration is 1%HF and concentration is 5%H 2o 2, make to compare O/A=1/1, wash, stir 4 minutes, 12 stage countercurrent washings; Add the NaOH of 2mol/L, maintain and compare O/A=1/1, stir and carry out back extraction in 30 minutes; Strip liquor filters, and supplementary NaOH recycles; Add pure water organic phase after organic phase back extraction, O/A=1/1, stir 10 ~ 15 minutes, clarification, aqueous phase can return washery slag when molten material filters; Then add 6mol/LHCl pickling, make to compare O/A=3 ~ 4/1, stir 10 ~ 15 minutes, clarification, spent acid returns molten material; Then add water and wash, compare O/A=1/10 ~ 15, stir 10 ~ 15 minutes, clarification, abandons aqueous phase;
2) back extraction thing NaOH back extraction got off filters, the supernatant liquor produced when filtrate can be used to precipitate below neutralization; Slag water is sized mixing, heating, add concentration be 10% HCL carry out excellent molten, control ph, at 1-2, stirs 0.5 hour, analytical solution zirconium content, in mass ratio phosphoric acid: zirconium=1:1 adds phosphoric acid, and steam heating is boiled, and stirs 1 hour, then adds HCl, reach 6N; Filter, check Sc in slag 2o 3content, waste;
3) filtrate is squeezed in extraction tank, adopt the organic phase of 38%P350,4%TBP and 58% kerosene composition, compare O/A=1/5,3 stage countercurrent extractions, mixing time 10-15 minute, join 6N washing lotion with pure water and HCl again, compare O/A=2:1, adverse current washes extraction 4 grades, with pure water back extraction scandium, back extraction is in a ratio of O/A=2/1, twice back extraction, mixing time 10 minutes;
4) by step 3) strip liquor that obtains squeezes into settling bath, is heated to 60-70 DEG C, in mass ratio oxalic acid: Scium trichloride=5.5:1 adds oxalic acid, limit edged stirs, and filters, and then dries filtering the oxalic acid scandium whizzer obtained, use pure water drip washing again, mother liquor and washing lotion alkali neutralize;
5) proceeding in watt alms bowl with the oxalic acid scandium of water wash, loading in retort furnace, first drying 2 hours at 200 DEG C, be then warming up to 700 DEG C, be incubated 4 ~ 6 hours; Cooling.
In above-mentioned tungsten slag valuable metal recovery process in, also comprise FeCl 3preparation process:
1) raffinate after extraction scandium is squeezed in groove, be heated to 95 DEG C, in hydrochloric acid, add MnO 2, make HCl content be 30g/L, keep solution limpid, be oxidized more than 50 minutes, check with or without ferrous ion, to during without ferrous ion, filtered while hot;
2) by the organic phase of 12 grades of 30%N235,20% sec-octyl alcohol and 50% kerosene group, O/A=1/1 ~ 5 are compared, ferric ion in extraction solution; Raffinate enters storage tank, then adds pure water, and carry out back extraction, O/A=3 ~ 5/1 is compared in back extraction, back extraction 5 grades; Use the organic phase back extraction 2 grades of 12 grades of 30%N235 of O/A=1:1 ~ 2,20% sec-octyl alcohol and 50% kerosene composition again, organic phase hcl acidifying, O/A=3.5/1, stirs 5 minutes; The outlet aqueous phase of second stage stripping process step is as the back extraction water of first step stripping process step;
3) strip liquor puts into enamel reaction still, heating condensing crystal.
In above-mentioned tungsten slag valuable metal recovery process in, also comprise Mn (OH) 2preparation process:
1) raffinate after extraction Fe is squeezed in groove, be heated to more than 85 DEG C, add liming and regulate pH to 5.5 ~ 6.0, boil 15 ~ 20 minutes, press filtration;
2) except step 18) filtrate that obtains is heated to more than 90 DEG C, under whipped state, slowly instill Na 2s solution, 3 ~ 5% add vulcanizing agent in mass ratio, stir 1 hour, the qualified rear press filtration of inspection heavy metal content, residue collection, and filtrate leaves standstill;
3), in the filtrate after removing heavy metals, at room temperature dripping mol ratio is the H of 1:3 2o 2and NH 3h 2the mixing solutions of O, stirs 30 minutes, generates hydrated manganese dioxide colloidal precipitation absorption except Si, leaves standstill, cross and filter insolubles;
4) filtrate after process is heated to 60 DEG C, stir and add NaOH solution, when reaction end is 6.5 ~ 7.0 to pH, be incubated 2 hours, press filtration obtains Mn (OH) 2.
In above-mentioned tungsten slag valuable metal recovery process in, also comprise CaCl 2preparation process:
To extract solution after Mn, and add in enamel reaction still, heating condensing crystal, dries throw out with whizzer, namely obtains CaCl 2.
Technique effect of the present invention is: the present invention has carried out synthetical recovery to tungsten, scandium, iron, manganese, tin, copper, tantalum niobium valuable element etc. in tungsten slag, reduces cost recovery, and the present invention also has good environmental benefit, advantage that cost is low.The present invention adopts acid dissolve tungsten slag to reclaim tungsten in acid condition and reclaims other elements simultaneously, and the technique of one alkali lixiviate tungsten fewer than existing wet method, technique is simpler.
Accompanying drawing explanation
Fig. 1 is process flow sheet of the present invention.
Embodiment
Below in conjunction with drawings and Examples, the present invention is further illustrated.
Embodiment 1
As shown in Figure 1, the present invention includes following steps:
One) preparation of tungsten
1) molten material
First the concentration adding calculated amount in cell body is 15 ~ 25% hydrochloric acid, be heated to 60 DEG C, black and white mixing tungsten slag is added at leisure under whipped state, after adding black and white mixing tungsten slag, add 15 ~ 25% hydrochloric acid of about 0.2 times of first time hydrochloric acid content again, after heated and boiled, add 20kg/t NaF, 10kg/t NaNO3; When pH value to about 1.0 time, then add 15 ~ 25% hydrochloric acid of 0.4 ~ 0.5 times amount of first time hydrochloric acid content, make S:L=1:4.5, boil when becoming reddish-brown to mud and stop heating, continue to stir, add water 2t/t, stirs 0.5h;
2) press filtration
When step 1) below the feed temperature to 80 prepared DEG C time, carries out filter press, after press filtration, add water washing, and air blast dries up; Filtrate is put into storage tank and is left standstill at least 12h, then essence is filtered;
3) alkali is molten
Toward groove in, first add the water of calculated amount, then by the 2nd) step press filtration wash after filter residue add, add 50 ~ 80kg/t NaOH simultaneously, add 20mg/t NaF, 20mg/t Na 3pO 4, L/S=1.5, heated and boiled 30 minutes, is incubated 1 hour, controls terminal pH10-11;
4) press filtration
When alkaline solution temperature is down to below 80 DEG C, carry out filter press, and with the washing slag 1 ~ 2 time of 30-40 DEG C; Slag rate 5-10%, Ta, Nb, Sn enrichment 8-16 doubly, can directly reclaim Ta, Nb, Sn valuable metal;
5) silica removal
Stirring and be 10% hydrochloric acid neutralization procedure 4 by concentration under boiling conditions) filtrate that produces to NaOH4-5g/L, add NH after boiling 30 minutes 4cl solution to pH value is 8 ~ 9, and sedimentation and filtration removes P, As, Si, filter cleaner;
6) ion-exchange
Adopt strong basicity gel-type anion resin by the filtrate ion-exchange absorption tungsten after molten for alkali silica removal; Wire velocity control 6-10cm/min during exchange, WO in liquid after inspection exit ion-exchange 3content; Leacheate is NaCl 4-8g/L or NaOH10g/L, drip washing linear velocity: 2-4cm/min, and stripping liquid is the NH of 3.5mol/L 4the NH of Cl or 2mol/L 3h 2o, desorb linear velocity: 24cm/min, stripping liquid consumption is 50% of resin volume;
7) W is extracted
By step 2) after the filtrate reduction of filtering of the essence that produces, the organic phase of 10 ~ 30%N235,10 ~ 20% sec-octyl alcohol and kerosene composition, squeezes in extraction tank by comparing O/A=1/20-30, stir 2 minutes, 4 stage countercurrent extractions, then add water by O/A=1/1, stir 3 minutes, 6 grades of washings; Then add the NaOH solution of the 2mol/L of heat by O/A=2/1, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then add water backwash by O/A=1/3 ~ 5, stir 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then add the hcl acidifying of 6mol/L by O/A=3 ~ 5/1, stir 3 ~ 5 minutes;
8) condensing crystal
Squeezed into by stripping liquid in enamel reaction still, calorify 80 ~ 85 DEG C of evaporative crystallizations, endpoint pH is 7 ~ 8, crystallization control rate 90% ~ 95%, crystal vacuum filtration, and with 30 ~ 40 DEG C of water washings, washing lotion and mother liquor return;
9) dry
By the crystal that washed 90 ~ 120 DEG C of dryings, obtain ammonium paratungstate.
Two) Sc 2o 3preparation
1) extract
By extraction W after more than liquid squeeze in extraction tank, add the organic phase of 12%P204,4%TBP and 84% kerosene composition, compare O/A=1/20, stir 3 minutes, 4 stage countercurrent extractions, control Scium trioxide content in raffinate and be less than 3mg/L, add that concentration is 15%HCl, concentration is 1%HF and concentration is 5%H 2o 2, make to compare O/A=1/1, wash, stir 4 minutes, 12 stage countercurrent washings; Add the NaOH of 2mol/L, maintain and compare O/A=1/1, stir and carry out back extraction in 30 minutes; Strip liquor filters, and supplementary NaOH recycles; Add pure water organic phase after organic phase back extraction, O/A=1/1, stir 10 ~ 15 minutes, clarification, aqueous phase can return washery slag when molten material filters; Then add the chlorohydric acid pickling of 6mol/L, make to compare O/A=3 ~ 4/1, stir 10 ~ 15 minutes, clarification, spent acid returns molten material; Then add water and wash, compare O/A=1/10 ~ 15, stir 10 ~ 15 minutes, clarification, abandons aqueous phase;
2) dissolve
Back extraction thing NaOH back extraction got off filters, the supernatant liquor produced when filtrate can be used to precipitate below neutralization; Slag water is sized mixing, heating, add concentration be 10% HCl carry out excellent molten, control ph, at 1-2, stirs 0.5 hour, analytical solution zirconium content, in mass ratio phosphoric acid: zirconium=1:1 adds phosphoric acid, and steam heating is boiled, and stirs 1 hour, then adds HCl, reach 6N; Filter, check Sc in slag 2o 3content, waste;
3) secondary extraction
Filtrate is squeezed in extraction tank, adopt the organic phase of 38%P350,4%TBP and 58% kerosene composition, compare O/A=1/5,3 stage countercurrent extractions, mixing time 10-15 minute, join 6mol/L washing lotion with pure water and HCl again, compare O/A=2:1, adverse current washes extraction 4 grades, with pure water back extraction scandium, back extraction is in a ratio of O/A=2/1, twice back extraction, mixing time 10 minutes;
4) oxalic acid precipitation
By step 3) strip liquor that obtains squeezes into settling bath, is heated to 60-70 DEG C, in mass ratio oxalic acid: Scium trichloride=5.5:1 adds oxalic acid, limit edged stirs, and filters, and then dries filtering the oxalic acid scandium whizzer obtained, use pure water drip washing again, mother liquor and washing lotion alkali neutralize;
5) calcination
Proceeding in watt alms bowl with the oxalic acid scandium of water wash, loading in retort furnace, first drying 2 hours at 200 DEG C, be then warming up to 700 DEG C, be incubated 4 ~ 6 hours; Namely cooling obtains Scium trioxide.
Three) FeCl 3preparation
1) oxidation of ferrous ion
Raffinate after extraction scandium is squeezed in groove, is heated to 95 DEG C, in hydrochloric acid, adds MnO 2, make HCl content be 30g/L, keep solution limpid, be oxidized more than 50 minutes, check with or without ferrous ion, to during without ferrous ion, filtered while hot;
2) ferric ion is extracted
By the organic phase of 12 grades of 30%N235,20% sec-octyl alcohol and 50% kerosene composition, compare O/A=1/1 ~ 5, ferric ion in extraction solution; Raffinate enters storage tank, then adds pure water, and carry out back extraction, O/A=3 ~ 5/1 is compared in back extraction, back extraction 5 grades; Use the organic phase back extraction 2 grades of 12 grades of 30%N235 of O/A=1:1 ~ 2,20% sec-octyl alcohol and 50% kerosene composition again, organic phase hcl acidifying, O/A=3.5/1, stirs 5 minutes; The outlet aqueous phase of second stage stripping process step is as the back extraction water of first step stripping process step.
3) concentrated
Strip liquor puts into enamel reaction still, and heating condensing crystal, namely obtains FeCl 3.
Four) Mn (OH) 2preparation
1) except Fe, As, Al, Si:
Raffinate after extraction Fe is squeezed in groove, is heated to more than 85 DEG C, add liming and regulate pH to 5.5 ~ 6.0, boil 15 ~ 20 minutes, press filtration;
2) removing heavy metals
Filtrate after Fe is heated to more than 90 DEG C, under whipped state, slowly instills Na 2s solution, 3 ~ 5% add vulcanizing agent in mass ratio, stir 1 hour, the qualified rear press filtration of inspection heavy metal content, residue collection, and filtrate leaves standstill;
3) degree of depth is except Si
In filtrate after removing heavy metals, at room temperature dripping mol ratio is the H of 1:3 2o 2and NH 3h 2the mixing solutions of O, stirs 30 minutes, generates hydrated manganese dioxide colloidal precipitation absorption except Si, leaves standstill, cross and filter insolubles;
4) precipitate
Filtrate after process is heated to 60 DEG C, and stir and add NaOH solution, when reaction end is 6.5 ~ 7.0 to pH, be incubated 2 hours, press filtration obtains Mn (OH) 2.
Five) CaCl 2preparation
To carry solution after Fe, Mn, and add in enamel reaction still, heating condensing crystal, dries with whizzer can obtain CaCl by throw out 2.
Embodiment 2
One) preparation of tungsten
1) molten material
Wash water after filtering with water or acidleach in cell body adds tungsten slag and sizes mixing under slow stirring state, and liquid-solid ratio is 1:1; Be steam heated to then 60 DEG C add 8% sulfuric acid according to the mass ratio 1:1.2 of tungsten slag and sulfuric acid, keep temperature of reaction 1.5 hours, then add 15% ~ 25%HCl of 0.25 times of sulfuric acid amount, after heated and boiled, then add 20kg/t NaF, 10kg/t NaNO 3; When pH value to about 1.0 time, put into 15 ~ 25% hydrochloric acid of 0.4 ~ 0.5 times of sulfuric acid amount, make S:L=1:4.5, boil when becoming reddish-brown to mud stop heating, continue stir, add water 2t/t, stir 0.5h; After feed temperature arrives 80 DEG C, by feed liquid under agitation, squeeze into filter and filter, then wash, dry up, filtrate is put into storage tank and is left standstill at least 12h, then essence is filtered.
2) press filtration
When step 1) below the feed temperature to 80 prepared DEG C time, can filter press be carried out, add water washing after press filtration, and air blast dry up; Filtrate is put into storage tank and is left standstill at least 12h, then essence is filtered;
3) alkali is molten
Toward groove in, first add the water of calculated amount, then by the 2nd) step press filtration wash after filter residue add, add 50 ~ 80kg/t NaOH simultaneously, add 20mg/t NaF, 20mg/t Na 3pO 4, L/S=1.5, heated and boiled 30 minutes, is incubated 1 hour, controls terminal pH10-11;
4) press filtration
When alkaline solution temperature is down to below 80 DEG C, carry out filter press, and with the washing slag 1 ~ 2 time of 30-40 DEG C; Slag rate 5-10%, Ta, Nb, Sn enrichment 8-16 doubly, can directly reclaim Ta, Nb, Sn valuable metal;
5) silica removal
Stirring and be 10% hydrochloric acid neutralization procedure 4 by concentration under boiling conditions) filtrate that produces to NaOH4-5gL, add NH after boiling 30 minutes 4cL solution to pH value is 8 ~ 9, and precipitation removes P, As, Si, filter cleaner;
6) ion-exchange
Adopt strong basicity gel-type anion resin by the filtrate ion-exchange absorption tungsten after molten for alkali silica removal; Wire velocity control 6-10cm/min during exchange, WO in liquid after inspection exit ion-exchange 3content; Leacheate is NaCl 4-8g/L or NaOH10g/L, drip washing linear velocity: 2-4cm/min, and stripping liquid is the NH of 3.5mol/L 4the NH of Cl or 2mol/L 3h 2o, desorb linear velocity: 24cm/min, stripping liquid consumption is 50% of resin volume;
7) W is extracted
By step 2) after the filtrate reduction of filtering of the essence that produces, the organic phase of 10 ~ 30%N235,10 ~ 20% sec-octyl alcohol and kerosene composition, squeezes in extraction tank by comparing O/A=1/20-30, stir 2 minutes, 4 stage countercurrent extractions, then add water by O/A=1/1, stir 3 minutes, 6 grades of washings; Then add the NaOH solution of the 2mol/L of heat by O/A=2/1, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then add water backwash by O/A=1/3 ~ 5, stir 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then add the hcl acidifying of 6mol/L by O/A=3 ~ 5/1, stir 3 ~ 5 minutes;
8) condensing crystal
Squeezed into by stripping liquid in enamel reaction still, calorify 80 ~ 85 DEG C of evaporative crystallizations, endpoint pH is 7 ~ 8, crystallization control rate 90% ~ 95%, crystal vacuum filtration, and with 30 ~ 40 DEG C of water washings, washing lotion and mother liquor return;
9) dry
By the crystal that washed 90 ~ 120 DEG C of dryings, obtain ammonium paratungstate.
Two) Sc 2o 3preparation
1) extract
By extraction W after more than liquid squeeze in extraction tank, add the organic phase of 12%P204,4%TBP and 84% kerosene composition, compare O/A=1/20, stir 3 minutes, 4 stage countercurrent extractions, control Scium trioxide content in raffinate and be less than 3mg/L, add that concentration is 15%HCl, concentration is 1%HF and concentration is 5%H 2o 2, make to compare O/A=1/1, wash, stir 4 minutes, 12 stage countercurrent washings; Add the NaOH of 2mol/L, maintain and compare O/A=1/1, stir and carry out back extraction in 30 minutes; Strip liquor filters, and supplementary NaOH recycles; Add pure water organic phase after organic phase back extraction, O/A=1/1, stir 10 ~ 15 minutes, clarification, aqueous phase can return washery slag when molten material filters; Then add the chlorohydric acid pickling of 6mol/L, make to compare O/A=3 ~ 4/1, stir 10 ~ 15 minutes, clarification, spent acid returns molten material; Then add water and wash, compare O/A=1/10 ~ 15, stir 10 ~ 15 minutes, clarification, abandons aqueous phase;
2) dissolve
The back extraction thing that NaOH back extraction got off filters, the supernatant liquor produced when precipitating below filtrate is used for neutralizing; Slag water is sized mixing, heating, add concentration be 10% HCl carry out excellent molten, control ph, at 1-2, stirs 0.5 hour, analytical solution zirconium content, in mass ratio phosphoric acid: zirconium=1:1 adds phosphoric acid, and steam heating is boiled, and stirs 1 hour, then adds HCL, reach 6mol/L; Filter, check Sc in slag 2o 3content, waste;
3) secondary extraction
Filtrate is squeezed in extraction tank, adopt the organic phase of 38%P350,4%TBP and 58% kerosene composition, compare O/A=1/5,3 stage countercurrent extractions, mixing time 10-15 minute, join 6mol/L washing lotion with pure water and HCl again, compare O/A=2:1, adverse current washes extraction 4 grades, with pure water back extraction scandium, back extraction is in a ratio of O/A=2/1, twice back extraction, mixing time 10 minutes;
4) oxalic acid precipitation
By step 3) strip liquor that obtains squeezes into settling bath, is heated to 60-70 DEG C, in mass ratio oxalic acid: Scium trichloride=5.5:1 adds oxalic acid, limit edged stirs, and filters, and then dries filtering the oxalic acid scandium whizzer obtained, use pure water drip washing again, mother liquor and washing lotion alkali neutralize;
5) calcination
Proceeding in watt alms bowl with the oxalic acid scandium of water wash, loading in retort furnace, first drying 2 hours at 200 DEG C, be then warming up to 700 DEG C, be incubated 4 ~ 6 hours; Namely cooling obtains Scium trioxide.
Three) FeCl 3preparation
1) oxidation of ferrous ion
Raffinate after extraction scandium is squeezed in groove, is heated to 95 DEG C, in hydrochloric acid, adds MnO 2, make HCl content be 30g/L, keep solution limpid, be oxidized more than 50 minutes, check with or without ferrous ion, to during without ferrous ion, filtered while hot;
2) ferric ion is extracted
By the organic phase of 12 grades of 30%N235,20% sec-octyl alcohol and 50% kerosene composition, compare O/A=1/1 ~ 5, ferric ion in extraction solution; Raffinate enters storage tank, then adds pure water, and carry out back extraction, O/A=3 ~ 5/1 is compared in back extraction, back extraction 5 grades; Use the organic phase back extraction 2 grades of 12 grades of 30%N235 of O/A=1:1 ~ 2,20% sec-octyl alcohol and 50% kerosene composition again, organic phase hcl acidifying, O/A=3.5/1, stirs 5 minutes; The outlet aqueous phase of second stage stripping process step is as the back extraction water of first step stripping process step.
3) concentrated
Strip liquor puts into enamel reaction still, and heating condensing crystal, namely obtains FeCl 3.
Four) Mn (OH) 2preparation
1) except Fe, As, Al, Si:
Raffinate after extraction Fe is squeezed in groove, is heated to more than 85 DEG C, add liming and regulate pH to 5.5 ~ 6.0, boil 15 ~ 20 minutes, press filtration;
2) removing heavy metals
Except filtrate after Fe is heated to more than 90 DEG C, under whipped state, slowly instill Na 2s solution, 3 ~ 5% add vulcanizing agent in mass ratio, stir 1 hour, the qualified rear press filtration of inspection heavy metal content, residue collection, and filtrate leaves standstill;
3) degree of depth is except Si
In filtrate after removing heavy metals, at room temperature dripping mol ratio is the H of 1:3 2o 2and NH 3h 2the mixing solutions of O, stirs 30 minutes, generates hydrated manganese dioxide colloidal precipitation absorption except Si, leaves standstill, cross and filter insolubles;
4) precipitate
Filtrate after process is heated to 60 DEG C, and stir and add NaOH solution, when reaction end is 6.5 ~ 7.0 to pH, be incubated 2 hours, press filtration obtains Mn (OH) 2.
Five) CaCl 2preparation
To carry solution after Fe, Mn, and add in enamel reaction still, heating condensing crystal, dries with whizzer can obtain CaCl by throw out 2.
Embodiment 3
Except one) 7 in the preparation process of tungsten) except extraction W step, all the other steps operate identical with embodiment 1, repeat no more.The extraction W step concrete operations of this embodiment are as follows:
By step 2) after the filtrate reduction of filtering of the essence that produces, squeezing in extraction tank by comparing O/A=1/20-30, stirring 2 minutes, adding concentration is 0.01% citric acid or stearic acid, and 4 stage countercurrent extractions, then add water by comparing O/A=1/1, stir 3 minutes, 6 grades of washings; Then by the NaOH solution of 2mol/L comparing O/A=2/1 and add heat, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then adding water backwash by comparing O/A=1/3 ~ 5, stirring 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then by the hcl acidifying comparing O/A=3 ~ 5/1 and add 6mol/L, stir 3 ~ 5 minutes.Add the process of the organic acid such as citric acid or stearic acid in extraction W process, can prevent from carrying secretly, be conducive to the recovery of precious metal.

Claims (6)

1. the recovery process of valuable metal in tungsten slag, comprises the steps:
1) concentration first adding calculated amount in cell body is 15 ~ 25% hydrochloric acid, be heated to 60 DEG C, black and white mixing tungsten slag is added at leisure under whipped state, after adding black and white mixing tungsten slag, add 15 ~ 25% hydrochloric acid of 0.2 times of first time hydrochloric acid content again, after heated and boiled, add 20kg/t NaF, 10kg/tNaNO 3; When pH value to about 1.0 time, then add 15 ~ 25% hydrochloric acid of first time hydrochloric acid content 0.4 ~ 0.5 times amount, make S:L=1:4.5, boil when becoming reddish-brown to mud and stop heating, continue to stir, add water 2t/t, stirs 0.5h;
2) when step 1) below the feed temperature to 80 prepared DEG C time, carries out filter press, after press filtration, water washing is added to filter residue, and air blast dries up; Filtrate is put into storage tank and is left standstill at least 12h, then essence is filtered;
3) first toward groove in, add water, then by the 2nd) step press filtration wash after filter residue add, add 50 ~ 80kg/t NaOH simultaneously, add 20kg/t NaF, 20kg/t Na 3pO 4, L/S=1.5, heated and boiled 30 minutes, be incubated 1 hour, control endpoint pH is 10-11;
4) when alkaline solution temperature is down to below 80 DEG C, filter press is carried out, and with the washing filter residue 1 ~ 2 time of 30-40 DEG C; Slag rate 5-10%, Ta, Nb, Sn enrichment 8-16 doubly, can directly reclaim Ta, Nb, Sn valuable metal;
5) stirring and be 10% hydrochloric acid neutralization procedure 4 by concentration under boiling conditions) filtrate that produces to NaOH 4-5g/L, add NH after boiling 30 minutes 4cl solution to pH value is 8 ~ 9, filter cleaner;
6) adopt strong basicity gel-type anion resin by the filtrate ion-exchange absorption tungsten after molten for alkali silica removal; Wire velocity control 6-10cm/min during exchange, WO3 content in liquid after inspection exit ion-exchange; Leacheate: NaCl 4-8g/L or NaOH10g/L, drip washing linear velocity 2-4cm/min, stripping liquid is: the NH of 3.5mol/L 4the NH of Cl or 2mol/L 3h 2o, desorb linear velocity: 24cm/min, stripping liquid consumption is 50% of resin volume;
7) by step 2) after the filtrate reduction of filtering of the essence that produces, the organic phase of 10 ~ 30%N235,10 ~ 20% sec-octyl alcohol and kerosene composition, squeeze in extraction tank by comparing O/A=1/20-30, stir 2 minutes, 4 stage countercurrent extractions, then add water by O/A=1/1, stir 3 minutes, 6 grades of washings; Then add the NaOH solution of the 2mol/L of heat by O/A=2/1, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then add water backwash by O/A=1/3 ~ 5, stir 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then add the hcl acidifying of 6mol/L by O/A=3 ~ 5/1, stir 3 ~ 5 minutes;
8) squeeze in enamel reaction still by stripping liquid, calorify 80 ~ 85 DEG C of evaporative crystallizations, endpoint pH is 7 ~ 8, crystallization control rate 90% ~ 95%, crystal vacuum filtration, and with 30 ~ 40 DEG C of water washings, washing lotion and mother liquor return absorption;
9) by the crystal that washed 90 ~ 120 DEG C of dryings, obtain ammonium paratungstate.
2. the recovery process of valuable metal in tungsten slag as claimed in claim 1, also comprises preparation Sc 2o 3step:
1) by extraction W after more than liquid squeeze in extraction tank, add the organic phase of 6 ~ 15%P204,3 ~ 6%TBP and kerosene composition, compare O/A=1/10 ~ 30, stir 3 minutes, 4 stage countercurrent extractions, control Scium trioxide content in raffinate and be less than 3mg/L, add that concentration is 15%HCl, concentration is 1%HF and concentration is 5%H 2o 2, make to compare O/A=1/1, wash, stir 4 minutes, 12 stage countercurrent washings; Add the NaOH of 2mol/L, heat 70 ~ 80 DEG C, compare O/A=1/1, stir and carry out back extraction in 30 minutes; Strip liquor filters, and supplementary NaOH recycles; Add 6mol/L HCl after organic phase back extraction and wash organic phase, compare O/A=3 ~ 4/1, stir 5 ~ 10 minutes, clarification, spent acid returns molten material; Then add water and wash, compare O/A=1/5 ~ 10, stir 10 ~ 15 minutes, clarification, abandons aqueous phase;
2) the back extraction thing that NaOH back extraction got off filters, the supernatant liquor produced when precipitating below filtrate is used for neutralizing; Slag water is sized mixing, heating, add concentration be 10% HCl carry out excellent molten, control ph, at 1-2, stirs 0.5 hour, analytical solution zirconium content, in mass ratio phosphoric acid: zirconium=1:1 adds phosphoric acid, and steam heating is boiled, and stirs 1 hour, then adds HCl, reach 6mol/L; Filter, check Sc in slag 2o 3content, waste;
3) filtrate is squeezed in extraction tank, adopt the organic phase of 30 ~ 45%P350,3 ~ 6%TBP and kerosene composition, compare O/A=1/5,3 stage countercurrent extractions, mixing time 5-10 minute, join 6mol/L washing lotion with pure water and HCl again, compare O/A=2:1, adverse current washes extraction 4 grades, with pure water back extraction scandium, back extraction is in a ratio of O/A=2/1, twice back extraction, mixing time 10 ~ 15 minutes;
4) by step 3) strip liquor that obtains squeezes into settling bath, is heated to 60-70 DEG C, in mass ratio oxalic acid: Scium trichloride=5.5:1 adds oxalic acid, limit edged stirs, and filters, and then dries filtering the oxalic acid scandium whizzer obtained, use pure water drip washing again, mother liquor and washing lotion alkali neutralize;
5) proceeding in watt alms bowl with the oxalic acid scandium of water wash, loading in retort furnace, first drying 2 hours at 200 DEG C, be then warming up to 750 DEG C, be incubated 2 ~ 4 hours; Namely cooling obtains Sc 2o 3.
3. the recovery process of valuable metal in tungsten slag as claimed in claim 2, also comprises FeCl 3preparation process:
1) raffinate after extraction scandium is squeezed in groove, be heated to 95 DEG C, add MnO 2, add HCl and make content be 30g/L, keep solution limpid, be oxidized more than 50 minutes, check with or without ferrous ion, to during without ferrous ion, filtered while hot;
2) by the organic phase of 12 grades of 30%N235,20% sec-octyl alcohol and 50% kerosene composition, O/A=1/1 ~ 5 are compared, ferric ion in extraction solution; Raffinate enters storage tank, then adds pure water, and carry out back extraction, O/A=3 ~ 5/1 is compared in back extraction, back extraction 5 grades; Use the organic phase back extraction 2 grades of 12 grades of 30%N235 of O/A=1:1 ~ 2,20% sec-octyl alcohol and 50% kerosene composition again, organic phase hcl acidifying, O/A=3.5/1, stirs 5 minutes; The outlet aqueous phase of second stage stripping process step is as the back extraction water of first step stripping process step;
3) strip liquor puts into enamel reaction still, and namely heating condensing crystal obtains FeCl 3.
4. the recovery process of valuable metal in tungsten slag as claimed in claim 3, also comprises Mn (OH) 2preparation process:
1) raffinate after extraction Fe is squeezed in groove, be heated to more than 85 DEG C, add liming and regulate pH to 5.5 ~ 6.0, boil 15 ~ 20 minutes, press filtration;
2) by step 1) filtrate that obtains is heated to more than 90 DEG C, under whipped state, slowly instill Na 2s solution, 3 ~ 5% add vulcanizing agent in mass ratio, stir 1 hour, the qualified rear press filtration of inspection heavy metal content, residue collection, and filtrate leaves standstill;
3), in the filtrate after removing heavy metals, at room temperature dripping mol ratio is the H of 1:3 2o 2and NH 3h 2the mixing solutions of O, stirs 30 minutes, generates hydrated manganese dioxide colloidal precipitation absorption except Si, leaves standstill, cross and filter insolubles;
4) filtrate after process is heated to 60 DEG C, stir and add NaOH solution, when reaction end is 6.5 ~ 7.0 to pH, be incubated 2 hours, press filtration obtains Mn (OH) 2.
5. the recovery process of valuable metal in tungsten slag as claimed in claim 4, also comprises CaCl 2preparation process:
To extract solution after Mn, and add in enamel reaction still, heating condensing crystal, dries throw out with whizzer, namely obtains CaCl 2.
6. the recovery process of valuable metal in tungsten slag as claimed in claim 1, described step 7) add citric acid or stearic acid process, its detailed process is as follows: by step 2) after the filtrate reduction of filtering of the essence that produces, squeezing in extraction tank by comparing O/A=1/20-30, stirring 2 minutes, adding concentration is 0.01% citric acid or stearic acid, 4 stage countercurrent extractions, then adding water by comparing O/A=1/1, stirring 3 minutes, 6 grades of washings; Then by the NaOH solution of 2mol/L comparing O/A=2/1 and add heat, stir 15 ~ 20 minutes; 3 grades of back extraction adverse currents, filter strip liquor; Then adding water backwash by comparing O/A=1/3 ~ 5, stirring 5 ~ 10 minutes, clarification, aqueous phase returns the molten use of alkali; Then by the hcl acidifying comparing O/A=3 ~ 5/1 and add 6mol/L, stir 3 ~ 5 minutes.
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CN1166789C (en) * 2002-04-10 2004-09-15 株洲硬质合金集团有限公司 United decomposition technology of tungstite, wolframite alkali decomposition
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CN102952951A (en) * 2011-08-22 2013-03-06 中国科学院过程工程研究所 Method for extracting tantalum and niobium and producing potassium fluosilicate from tungsten smelting slag

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